Method for recovering metal values from refractory sulfide ore

ABSTRACT

A method for recovering metal values from refractory sulfide ores is provided. The method includes the steps of separating clays and fines from crushed refractory sulfide ore, forming a heap from the refractory sulfide ore, producing a concentrate of refractory sulfide minerals from the separated fines and adding the concentrate to the heap, bioleaching the heap to thereby oxidize iron sulfides contained therein, and hydrometallurgically treating the bioleached ore to recover metal values contained therein.

PRIORITY INFORMATION

This is a continuation of application Ser. No. 12/286,653, filed Sep.30, 2008, now U.S. Pat. No. 7,641,714, which is a continuation of Ser.No. 11/619,186, filed Jan. 2, 2007, now U.S. Pat. No. 7,429,286, whichis a continuation of application Ser. No. 10/720,555, filed Nov. 24,2003, now U.S. Pat. No. 7,156,894, which is a continuation ofapplication Ser. No. 10/146,393, filed May 14, 2002, now U.S. Pat. No.6,652,622, which is a continuation of application Ser. No. 09/709,765,filed Nov. 10, 2000, now U.S. Pat. No. 6,387,155, which is acontinuation of application Ser. No. 08/950,279, filed Oct. 14, 1997,now U.S. Pat. No. 6,146,444, which is a continuation of application Ser.No. 08/476,444, filed Jun. 7, 1995, now U.S. Pat. No. 5,676,733, whichis a continuation in part of application Ser. No. 08/343,888, filed Nov.16, 1994, now U.S. Pat. No. 5,573,575, which is a continuation in partof application Ser. No. 08/161,742, filed Dec. 3, 1993, now U.S. Pat.No. 5,431,717.

TECHNICAL FIELD

The present invention relates to the recovery of metal values fromrefractory sulfide and refractory carbonaceous sulfide ores.

BACKGROUND ART

Gold is one of the rarest metals on earth. Gold ores can be categorizedinto two types: free milling and refractory. Free milling ores are thosethat can be processed by simple gravity techniques or directcyanidation. Refractory ores, on the other hand, are not amenable toconventional cyanidation treatment. Such ores are often refractorybecause of their excessive content of metallic sulfides (e.g., pyrite)and/or organic carbonaceous matter.

A large number of refractory ores consist of ores with a precious metalsuch as gold occluded in iron sulfide particles. The iron sulfideparticles consist principally of pyrite and arsenopyrite. Precious metalvalues are frequently occluded within the sulfide mineral. For example,gold often occurs as finely disseminated sub-microscopic particleswithin a refractory sulfide host of pyrite or arsenopyrite. If the goldremains occluded within the sulfide host, even after grinding, then thesulfides must be oxidized to liberate the encapsulated precious metalvalues and make them amenable to a leaching agent (or lixiviant).

A number of processes for oxidizing the sulfide minerals to liberate theprecious metal values are well known in the art. One known method ofoxidizing the metal sulfides in the ore is to use bacteria, such asThiobacillus ferrooxidans, Sulfolobus, Acidianus species andfacultative-thermophilic bacteria in a microbial pretreatment. Theforegoing microorganisms oxidize the iron sulfide particles to cause thesolubilization of iron as ferric iron, and sulfide, as sulfate ion.

If the refractory ore being processed is a carbonaceous sulfide ore,then additional process steps may be required following microbialpretreatment to prevent preg-robbing of the aurocyanide complex or otherprecious metal-lixiviant complexes by the native carbonaceous matterupon treatment with a lixiviant.

As used herein, sulfide ore or refractory sulfide ore will be understoodto also encompass refractory carbonaceous sulfide ores.

A known method of bioleaching carbonaceous sulfide ores is disclosed inU.S. Pat. No. 4,729,788, issued Mar. 8, 1988, which is herebyincorporated by reference. According to the disclosed process,thermophilic bacteria, such as Sulfolobus and facultative-thermophilicbacteria, are used to oxidize the sulfide constituents of the ore. Thebioleached ore is then treated with a blanking agent to inhibit thepreg-robbing propensity of the carbonaceous component of the ore. Theprecious metals are then extracted from the ore using a conventionallixiviant of cyanide or thiourea.

Another known method of bioleaching carbonaceous sulfide ores isdisclosed in U.S. Pat. No. 5,127,942, issued Jul. 7, 1992, which ishereby incorporated by reference. According to this method, the ore issubjected to an oxidative bioleach to oxidize the sulfide component ofthe ore and liberate the precious metal values. The ore is theninoculated with a bacterial consortium in the presence of nutrientstherefor to promote the growth of the bacterial consortium, thebacterial consortium being characterized by the property of deactivatingthe preg-robbing propensity of the carbonaceous matter in the ore. Inother words, the bacterial consortium functions as a biological blankingagent. Following treatment with the microbial consortium capable ofdeactivating the precious-metal-adsorbing carbon, the ore is thenleached with an appropriate lixiviant to cause the dissolution of theprecious metal in the ore.

Problems exist, however, with employing bioleaching processes in a heapleaching environment. These include nutrient access, air access, andcarbon dioxide access for making the process more efficient and thus anattractive treatment option. Moreover, for biooxidation, the inductiontimes concerning biooxidants, the growth cycles, viability of thebacteria and the like are important considerations because the variablessuch as accessibility, particle size, settling, compaction and the likeare economically irreversible once a heap has been constructed. As aresult, heaps cannot be repaired once formed, except on a limited basis.

Ores that have a high clay and/or fines content are especiallyproblematic when processing in a heap leaching or heap biooxidationprocess. The reason for this is that the clay and/or fines can migratethrough the heap and plug channels of air and liquid flow, resulting inpuddling; channeling; nutrient-, carbon dioxide-, or oxygen-starving;uneven biooxidant distribution, and the like. As a result, large areasof the heap may be blinded off and ineffectively leached. This is acommon problem in cyanide leaching and has lead to processes of particleagglomeration with cement for high pH cyanide leaching and with polymersfor low pH bioleaching. Polymer agglomerate aids may also be used inhigh pH environments, which are customarily used for leaching theprecious metals, following oxidative bioleaching of the iron sulfides inthe ore.

Biooxidation of refractory sulfide ores is especially sensitive toblocked percolation channels by loose clay and fine material because thebacteria need large amounts of air or oxygen to grow and biooxidize theiron sulfide particles in the ore. Air flow is also important todissipate heat generated by the exothermic biooxidation reaction,because excessive heat can kill the growing bacteria in a large, poorlyventilated heap.

Ores that are low in sulfide or pyrite, or ores that are high in acidconsuming materials such as calcium carbonate or other carbonates, mayalso be problematic when processing in a heap biooxidation. The reasonfor this is that the acid generated by these low pyrite ores isinsufficient to maintain a low pH and high iron concentrate needed forbacteria growth.

A need exists, therefore, for a heap bioleaching technique that can beused to biooxidize precious metal bearing refractory sulfide ores andwhich provides improved air and fluid flow within the heap. In addition,a need exists for a heap bioleaching process in which ores that are lowin sulfide minerals, or ores that are high in acid consuming materialssuch as calcium carbonate, may be processed.

A need also exists for a biooxidation process which can be used toliberate occluded precious metals in concentrates of refractory sulfideminerals. Mill processes that can be used for oxidizing suchconcentrates include bioleaching in a stirred bioreactor, pressureoxidation in an autoclave, and roasting. These mill processes oxidizethe sulfide minerals in the concentrate relatively quickly, therebyliberating the entrapped precious metals. However, unless theconcentrate has a high concentration of gold, it does not economicallyjustify the capital expense or high operating costs associated withthese processes. And, while a mill bioleaching process is the leastexpensive mill process in terms of both the initial capital costs andits operating costs, it still does not justify processing concentrateshaving less than about 0.5 oz. of gold per ton of concentrate, whichtypically requires an ore having a concentration greater than about 0.07oz. of gold per ton. Therefore, a need also exists for a process thatcan be used to biooxidize concentrates of precious metal bearingrefractory sulfide minerals at a rate comparable to a stirred tankbioreactor, but that has capital and operating costs more comparable tothat of a heap bioleaching process.

In addition to concentrates of precious metal bearing sulfide minerals,there are many sulfide ores that contain metal sulfide minerals that canpotentially be treated using a biooxidation process. For example, manycopper ores contain copper sulfide minerals. Other examples include zincores, nickel ores, and uranium ores. Biooxidation could be used to causethe dissolution of metal values such as copper, zinc, nickel and uraniumfrom concentrates of these ores. The dissolved metal values could thenbe recovered using known solvent extraction techniques, ironcementation, and precipitation. However, due to the sheer volume of thesulfide concentrate formed from sulfide ores, a stirred bioreactor wouldbe prohibitively expensive, and standard heap operations would simplytake too long to make it economically feasible to recover the desiredmetal values. A need also exists, therefore, for an economical processfor biooxidizing concentrates of metal sulfide minerals produced fromsulfide ores to thereby cause the dissolution of the metal values sothat they may be subsequently recovered from the bioleachate solution.

SUMMARY

It is an object of one aspect of the present invention to provide a heapbioleaching process of the type described above, wherein the refractorysulfide ore is rendered more susceptible to biooxidation, therebyproviding improved recovery of the precious metal values containedwithin the ore. The method of the present invention achieves this objectby removing the clays and/or fines from the refractory sulfide ore afterit is crushed to a size appropriate for a heap leaching process. Theheap may then be formed without concern of the air and liquid flowchannels in the heap becoming plugged. Further, if the separated clayand/or fine material has a sufficiently high precious metal content, itmay be separately treated to recover the precious metal values containedtherein.

In another aspect of the present invention, a process for recoveringprecious metal values from concentrates of precious metal bearingrefractory sulfide minerals is provided. The process comprises the stepsof (a) distributing a concentrate of refractory sulfide minerals on topof a heap of support material; (b.) biooxidizing the concentrate ofrefractory sulfide minerals; (c.) leaching precious metal values fromthe biooxidized refractory sulfide minerals with a lixiviant; and (d.)recovering precious metal values from the lixiviant. An advantage ofthis process is that the rate at which the sulfide minerals biooxidizeis much higher than would be observed in a traditional heap bioleachingoperation. Despite this high rate of biooxidation, however, the initialcapital costs and operating costs for the disclosed process are lowerthan that associated with a mill type biooxidation process.

Gold is the preferred precious metal recovered using the processaccording to the present aspect of the invention. However, otherprecious metals can also be recovered, including silver and platinum.The support material is preferably selected from the group consisting oflava rock, gravel, and coarsely ground ore. Lava rock is particularlypreferred due to its high surface area. As those skilled in the art willimmediately recognize, a number of lixiviants can be used in conjunctionwith the present process, however, thiourea and cyanide are thepreferred, cyanide being a particularly preferred lixiviant.

In another aspect of the present invention a process is provided forrecovering metal values from sulfide ores. Such ores include, by way ofexample, chalcopyrite, sphalorite, nickel sulfide ores, and uraniumsulfide ores. The process according to this aspect of the inventioncomprises (a.) forming a concentrate of metal sulfide minerals; (b.)spreading the concentrate on top of a heap of support material; (c.)biooxidizing the concentrate; and (d.) recovering metal values from thesolution used to biooxidize the metal sulfide minerals. Due to the factthat this process uses a heap of support material for the bioreactor,its capital and operating costs are less than that of a mill bioleachingoperation. However, due to the good air flow in the heap, thebiooxidation rate of the sulfide minerals is quite high and can approachthat of what would be observed in a mill type operation.

Depending on the sulfide ore from which the concentrate is obtained, themetal values that can be recovered from the process according to thepresent aspect of the invention include copper, zinc, nickel anduranium. The support material used in the present process is preferablyselected from the group consisting of lava rock, gravel, and coarselyground rock. Lava rock is particularly preferred due to its high surfacearea.

The above and other objects, features and advantages will becomeapparent to those skilled in the art from the following description ofthe preferred embodiments.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a schematic of a process flow sheet according to a preferredembodiment of the present invention;

FIG. 2 is a graph illustrating the percent iron leached over time forvarious size fractions of ore;

FIG. 3 is a graph illustrating the percent iron leached per day as afunction of time for various size fractions of ore;

FIG. 4 is a graph illustrating the percent gold recovered from a pyriteconcentrate milled to −200 mesh as a function of its percentbiooxidation;

FIG. 5 is a graph illustrating the change in Eh of a column of +¼ inchore as a function of time;

FIG. 6 is a graph illustrating the change in pH as a function of timefor a column of +¼ inch ore;

FIG. 7 is a graph illustrating the change in iron concentration in theeffluent of a column of +¼ inch ore as a function of time; and

FIG. 8 is a graph illustrating the biooxidation rate for a concentrateon lava rock process according to the present invention versus a stirredtank type process.

DETAILED DESCRIPTION

According to one aspect of the present invention, refractory sulfideores can be rendered more susceptible to biooxidation in a heap leachingprocess. This is accomplished by separating the clay and/or finematerials from the refractory sulfide ore after it has been crushed to asize appropriate for heap leaching. In the present embodiment the methodof removal is wet size screening. It will be readily apparent to thoseskilled in the art, however, that any other method for separating theclay and/or fine material from the refractory ore may be used. Forexample, dry screening and cyclone classifying are well known to thoseskilled in the art.

By removing the fines and clays from the refractory sulfide ore, the airand liquid flow through the heap is improved. This will reduce the timerequired to sufficiently biooxidize the iron sulfide particles in theore to liberate the precious metal values and make them amenable tosubsequent lixiviation with cyanide or thiourea, preferably cyanide. Inaddition to faster biooxidation, in a well ventilated heap, having goodfluid flow, it becomes more feasible to change the pH from an acidic pHof 1.0 to 2.0 that is best for biooxidation to a basic pH of 10.0 ormore needed for cyanide leaching without remaking or restacking theheap.

The refractory sulfide ore is preferably crushed to a target maximumsize in the range of approximately ¼ to 1 inch. Appropriate targetmaximum particle sizes include ¼, ⅜, ½, and 1 inch. If the ore will passany of these target particle sizes, it should be amenable to heapleaching. The smaller the particle size, however, the greater thesurface area of the sulfide particles in the ore and, of course, thefaster the sulfide particles will be biooxidized. Increased recovery ofthe precious metal values should also result. This, however, must beweighed against the additional cost of crushing the ore to a smallerparticle size. The additional amount of precious metal recovered may notjustify the added cost.

Of course if the refractory sulfide ore body being treated is already anappropriate size for heap leaching, no additional crushing is required.

Fines are naturally produced during the crushing process. The size ofthe fines and clays removed from the crushed ore should be about minus60 mesh as a minimum upper limit to about minus ⅛ inch as a maximumupper limit. After the clay and fines are separated from the bulk of theore, a heap is formed with the ore. The heap may then be treated with astandard bioleaching process to oxidize the iron sulfide particles inthe ore and liberate the occluded precious metal values, which arepreferably gold. Because the majority of the clay and fine materialshave been removed, obstruction of the air and liquid flow channels bythese materials is no longer a concern, thereby improving percolationleaching of the ore.

After biooxidation, the precious metal in the pretreated ore can beextracted using a conventional lixiviant such as cyanide or thiourea,preferably cyanide. Of course, however, as a person of ordinary skill inthe art would recognize, if the refractory sulfide ore is alsorefractory due to carbonaceous matter contained in the ore, additionalprocessing steps must be employed to reduce the preg-robbing propensityof the ore prior to lixiviation. A number of such processes are wellknown in the art.

For example, the methods used in U.S. Pat. No. 4,729,788 and U.S. Pat.No. 5,127,942, both of which have already been incorporated herein byreference, can be used. Further, the microbial process for treatingcarbonaceous ores disclosed in U.S. Pat. No. 5,162,105, issued Nov. 10,1992, hereby incorporated by reference, can also be used.

The fine material that has been separated may contain large amounts ofprecious metal values. Indeed the economic value of these metal valuesmay be sufficiently high to justify further processing of thesematerials to recover the additional metal values. In a particularlypreferred embodiment of the present invention, the separated finematerial is further processed to recover at least a portion of theprecious metal values contained therein.

To recover the precious metal values from the fine material, the finematerial is preferably treated in a mill process to remove the ironsulfide particles from the clay and sand particles. The reason for thisis that, as discussed above, precious metal values, especially gold,often occur as finely disseminated microscopic particles within the ironsulfide particles. These fine sulfide particles, therefore, frequentlycontain a significant portion of the overall precious metal values.Further, because a relatively high percentage of the precious metalvalues in the ore are associated with this fraction of the ore, they canbe economically treated in a mill process.

As will be recognized by those skilled in the art, a variety of methodscan be used to separate the iron sulfide particles from the remainder ofthe fines. These methods include, by way of example only, gravityseparation and flotation. If desired, the iron sulfide particles can besubjected to additional grinding before flotation. Gravity separationtechniques that can be used include shaker tables, hydrocyclones, andspiral classifiers.

The iron sulfide concentrate, if refractory, is preferably bioleachedwith bacteria in a tank or mill process to liberate the occludedprecious metal values. Alternatively, the sulfide concentrate can beadded back to the heap to allow for a slower heap biooxidation process.However, because these particles are typically larger and morehydrophobic than clay particles, they tend to stick more readily to thelarger particles in the heap, and, thus, the problem of obstructedpercolation channels is not encountered. The iron sulfide concentratecan also be treated by a variety of other methods well known in the artsuch as roasting, pressure oxidation, and chemical oxidation. Becausethe concentration of gold or other precious metal values is relativelyhigh in this ore fraction and its overall volume small, all of thesemill processes may be economically utilized.

If the iron sulfide concentrate is only partially refractory, then itcan be directly leached with a lixiviant such as cyanide to remove thenonrefractory gold. The tail from this leaching process could then bewashed free of cyanide and added to the heap for biooxidation to releasethe remaining refractory gold or other precious metal values.

The fine material removed from the refractory sulfide ore by sizeseparation, and which has also had the iron sulfide particles removedfrom it, may still contain economic values of gold or other preciousmetals. Further, this fine material is likely to be less refractory thanother iron sulfide material if the size has lead to oxidation.Therefore, agglomeration of this material with cement, or otheragglomeration aids that can be used at a high pH, may provide goodrecoveries if leached with cyanide directly.

The fine material may have sufficient gold value in the case of highgrade ore to merit a mill leaching process such as carbon-in-pulp orcounter current decantation.

A more recently preferred embodiment of the present invention is nowdescribed in connection with the process flow sheet illustrated in FIG.1.

As can be seen from referring to FIG. 1, a precious metal bearingrefractory sulfide ore is preferably crushed to a target maximum size inthe range of approximately ¼ to 1 inch at crushing station 10.Preferably the ore is crushed to a target maximum particle size of ¼, ⅜,½, or ¾ inch. Of course, if the refractory sulfide ore body beingtreated is already of an appropriate size for heap leaching, noadditional crushing is required.

As in the present embodiment, the precious metal to be recovered fromthe ore is typically gold. However, as those skilled in the art willreadily recognize, the method according to the present invention isequally applicable to the recovery of other precious metals, includingsilver and platinum from refractory sulfide ores.

After the gold bearing refractory sulfide ore is crushed to theappropriate size, the fines in the ore are separated from the crushedore at separation station 12. Preferably the fines are separated using awet or dry screening process. To ensure good air and liquid flow in theheap, fines smaller than about 10 to 30 mesh (Tyler mesh series) shouldbe separated out at separation station 12. The coarse fraction of theore 14, that is the ore greater than about 10 to 30 mesh, will typicallycontain approximately 50% or more of the gold values in the entire oreand comprise about 50 to 75% of the weight of the ore. The fines 16 thathave been separated out will typically contain approximately 30 to 50%of the gold values and comprise approximately 25 to 50% of the weight ofthe initial ore.

Because of the significant gold values typically contained in the fines16, the fines are further processed to recover at least a portion of theprecious metal values contained therein. This is preferably accomplishedby producing a concentrate 20 of refractory pyrite minerals from thefines 16 in the pyrite concentration circuit 22. Pyrite concentrate 20will typically comprise about 5 to 10% of the initial weight of the oreand about 15 to 30% of its gold values.

If the ore contains refractory arsenopyrite minerals, then refractorypyrite concentrate 20 will also contain these minerals.

Because, as a general rule, the pyrite particles in the pyriteconcentrate 20 are larger and more hydrophobic than the clay particlesfound in fines 16, the pyrite concentrate 20 can be combined with thecoarse fraction of the ore 14 during heap construction withoutsignificantly impeding fluid and air flow within the heap duringbioleaching. This is because the pyrite particles in pyrite concentrate20 will tend to stick to the larger particles in the coarse fraction ofthe ore, rather than migrating through the heap and causing blocked flowchannels. Pyrite concentrate 20, may also be added to the top of theheap before or after the biooxidation process is already in progress.

The bacterial oxidation of pyrite generates ferrous sulfate and sulfuricacid in the net reaction summarized by Equation (1). This net reactioncan be broken into two distinct reactions, Equations (2) and (3), whereEquation (2) is the aerobic reaction catalyzed by bacterial activity andEquation (3) is the anaerobic reaction occurring at the surface of thesulfide mineral. Equation (4) is a similar anaerobic reaction occurringat the surface of arsenopyrite minerals.FeS₂+7/2O₂+H₂O=FeSO₄+H₂SO₄  (1)14FeSO₄+7H₂SO₄+7/2O₂=7H₂O+7Fe₂(SO₄)₃  (2)7Fe₂(SO₄)₃+FeS₂+8H₂O=15FeSO₄+8H₂SO₄  (3)13Fe₂(SO₄)₃+2FeAsS+16H₂O=20FeSO₄+2H₃AsO₄+13H₂SO₄  (4)

An advantage of adding pyrite concentrate 20 to heap 18 is that thisfine milled pyrite is more readily oxidized than the pyrite mineralparticles found in coarse ore 14; thus, the acid produced from theoxidation of the pyrite concentrate can be used to help lower the pH ofthe coarse ore in the heap more quickly. This is especially valuablewhen dealing with ores that are high in acid consuming materials such ascalcium carbonate or other carbonates. Further, by adding the pyriteconcentrate to the top of heap 18, ferric ions produced during itsbiooxidation will migrate to the lower part of the heap where bacterialgrowth may be inhibited due to toxins, which have not been washed fromthe ore early in the biooxidation process, or due to the lack of oxygen.As a result, biooxidation of the pyrite minerals in the lower part ofthe heap will proceed even if bacterial growth is not favored in thisregion.

There is also an advantage to adding pyrite concentrate 20 to a heap 18that has been undergoing biooxidation for a long period of time. In thelater stages of biooxidation most of the exposed and reactive sulfideswill have already been oxidized, resulting in a slow down in the rate ofbiooxidation. This slow down in the rate of biooxidation can lead to adrop in iron levels and an increase in pH within heap 18. Addition of areactive sulfide concentrate can restart an active biooxidation processthat can increase indirect chemical leaching of imbedded sulfideminerals due to the high ferric levels produced from the biooxidation ofthe sulfide concentrate.

The preferred methods of producing pyrite concentrate 20 are explainedin detail below in connection with pyrite concentration circuit 22.

After heap 18 is constructed, it may be pretreated using a standard heapbiooxidation process to oxidize the iron sulfide particles in the oreand liberate the occluded precious metal values. And, because themajority of the clay and fine materials have been removed, obstructionof the air and liquid flow channels by these materials is significantlyreduced, resulting in improved percolation leaching of the ore.

If the bioleachate solution is recycled during the biooxidation process,the biooxidation rate can be improved by using the method of solutionmanagement disclosed in the U.S. patent application Ser. No. 08/329,002,entitled “Method For Improving The Heap Biooxidation Rate Of RefractorySulfide Ore Particles That Are Biooxidized Using Recycled BioleachateSolution,” which was filed Oct. 25, 1994, by William J. Kohr, ChrisJohannson, John Shield, and Vandy Shrader, the text of which isincorporated herein by reference as if fully set forth.

Referring again to FIG. 1, pyrite concentration circuit 22 is nowdescribed. Three preferred methods of producing pyrite concentrate 20are illustrated within pyrite concentration circuit 22. These methodsmay be used in combination or in the alternative.

The fines 16 will typically comprise very fine clay particles, which aretypically less than 5 to 20 μm; sand particles; and refractory sulfideparticles. The clay particles are very small and very hydrophilic incomparison to the sand and refractory sulfide particles, making themparticularly deleterious to heap bioleaching processes, because theytend to migrate through the heap and plug flow channels as they swellfrom the absorption of water. The clay particles are, therefore,preferably removed from the fines 16 so that a concentrate of refractorysulfide particles can be produced that can be safely added to heap 18with minimal obstruction of the flow channels in the heap. Thus, asillustrated in FIG. 1, the first step in each of the preferred methodsof producing pyrite concentrate 20 is the removal of the clay particlesfrom the fines using clay removal system 24, which is preferably ahydrocyclone or a settling tank. Of course, however, if the ore is a lowclay bearing ore, this step may be omitted.

The set point for the maximum size particle removed in clay removalsystem 24 will depend on the distribution of clay particle sizes withinfines 16. If the set point for the clay removal system is set at lessthan about 10 μm, a settling tank is the preferred removal method ofseparation because hydrocyclones cannot currently make efficientseparations between particle sizes of less than about 10 μm.

In a high clay ore, clay material 26 separated from the fines 16 willtypically comprise about 10% of the initial weight of the ore and about5 to 10% of its gold values. Further, because of its low refractorynature, clay material 26 may be further processed to recover the goldvalues it contains using a traditional cyanide mill leaching processsuch as counter current decantation or carbon-in-pulp. Before processingclay material 26 in one of these traditional cyanide mill leachingprocesses, however, the pulp density of the clay material should beincreased using a thickener 28 until a pulp density of about 30 to 40%is achieved.

After the clays have been removed from the fines 16, the refractorypyrite particles are also separated out to form refractory pyriteconcentrate 20, which can be added to heap 18 as explained above. Therefractory pyrite particles are preferably separated from clay depletedfines 30 using flotation or a gravity separation technique.

Three preferred methods for separating the refractory sulfide particlesfrom the clay depleted fines 30 are now described. The first methodentails fine grinding the clay depleted fines 16 until a particle sizeof less than about −200 mesh is achieved. This is preferablyaccomplished in ball mill 34. The refractory pyrite materials are thenremoved from the material 30 using a flotation cell 36 with a xanthatecollector. The floated pyrite material from flotation cell 36 forms thepyrite concentrate 20.

A second method of producing pyrite concentrate 20 from material 30comprises separating material 30 into two fractions using a hydrocyclone38: the first, comprising −200 mesh material 40, and the secondcomprising coarse sand particles, which are greater than about 200 mesh,and heavier pyrite particles. The material which is less than 200 meshis further treated in xanthate flotation cell 36 to remove refractorysulfides. The floated refractory sulfides and the coarse sand particlesand heavier pyrite are then recombined to form pyrite concentrate 20.This method differs from the first pyrite concentration method in thatinstead of crushing all of material 30 to less than −200 mesh, the sandparticles greater than 200 mesh and the heavier pyrite minerals inmaterial 30 are simply separated from material 30 and then added to thefloated pyrite from the −200 mesh material 40.

The third method of producing pyrite concentrate 20 from clay depletedfines 30 comprises using a gravity technique such as a spiral classifieror shaker table to remove the heavier sulfide minerals from theremainder of material 30.

The tail material 32, which remains after the refractory sulfidefraction has been removed from the clay depleted fines material 30,comprises approximately 20 to 30% of the initial weight of the ore andabout 5 to 10% of its gold, approximately 85% of which is recoverable ina traditional cyanide mill leaching process such as counter currentdecantation or carbon-in-pulp. Thus, tail material 32 is not veryrefractory and may be treated with clay material 26 in a traditionalmill cyanide leaching process to help improve the overall recovery ofthe process.

After heap 18 is biooxidized, the precious metal in the pretreated orecan be extracted using a conventional lixiviant such as cyanide orthiourea, preferably cyanide. Of course, however, as a person ofordinary skill in the art would recognize, if the refractory sulfide oreis also refractory due to carbonaceous matter contained in the ore,additional processing steps must be employed to reduce the preg-robbingpropensity of the ore prior to lixiviation as explained above.

EXAMPLE 1

A sample of 16 kg of refractory sulfide ore with approximately 0.04oz/ton of gold and 3.5% of sulfide sulphur was crushed to −¼ inch. Theore sample was then separated by wet screening into a +⅛ to −¼ inch, a+30 mesh to −⅛ inch, and a −30 mesh material fractions. The −30 meshmaterial was further separated into a pyrite fraction, a sand fraction,and a clay fraction by gravity separation. The sand fraction was furtherprocessed by fine grinding in a ball mill for about one hour. Thismaterial was then floated with xanthate as a collector.

Each fraction was then dried and weighed and analyzed for gold. The +⅛to −¼ inch material represented 51% of the weight and 18% of the gold at0.48 ppm Au. The +30 mesh to −⅛ inch material represented 28% of theweight and 32% of the gold at 1.47 ppm Au. The total pyrite, whichincluded both the gravity separated pyrite and the pyrite concentratefrom the flotation of the sand, represented 4.7% of the weight and 35%of the gold at 9.8 ppm Au. The remaining sand flotation tail and claymaterial represented 16% of the weight and 14.6% of the gold at about1.2 ppm Au.

The +⅛ to −¼ inch material and the +30 mesh to −⅛ inch material werecombined according to their weight percentages. The combined materialwas adjusted to a pH of 2.0 with 10% sulfuric acid at 30 ml/kg. The onemixture was then poured into a column and aerated from the bottom withat least 15 l of air/min/m² and liquid dilute basal solutions of(NH₄)₂SO₄ 0.04 g/l MgSO₄.7H₂O at 0.04 g/l and KH₂PO₄ at 0.004 g/l wereadded to the top at about 15 ml/hour. Thiobacillus ferrooxidans bacteriawere added to the top of the column and washed into the column with theliquid flow. This procedure allowed for good air flow and liquid flowand also migration of bacteria through the column. After about one monththe effluent from the column showed good bioleaching of iron at about0.1% per day.

EXAMPLE 2

A second sample of ore from the same mine as in Example 1 was crushed to−⅜ inches. Four 23 Kg splits of this sample were combined and wetscreened into a +¼ inch, a +⅛ to −¼ inch, a +10 mesh to −⅛ inch, a +16mesh to −10 mesh, a +30 to −16 mesh, a +60 to −30 mesh, and a −60 meshfraction. The +60 to −30 mesh and the −60 mesh fraction were used toevaluate a number of gravity separations to make a pyrite fraction asand fraction and a clay fraction. The dry weights of each size fractionwere used to calculate the weight percentage of the size fraction. Eachsize fraction was also analyzed for the amount of gold, iron and goldextraction by traditional cyanide leaching (see Table 1).

The five size fractions larger than 30 mesh were put into individualcolumns for biooxidation. Bacteria and nutrients were added as inExample 1 and air was blown in from the bottom or top of the column. Theprogress of the biooxidation was monitored by measuring the amount ofiron leached from the ore by using atomic absorption analysis of thenutrient solution passing through the column. The approximate totalamount of iron in each size fraction of the ore was calculated from theweight of the size fraction and an iron analysis of a representativesample of the ore. The percent iron leached and the average percent ironleached each day are plotted against time for all five size fractions inFIGS. 2 and 3, respectively.

TABLE 1 Ore Size Fraction Analysis GRAVITY SEPARATION Au Au % BIOOX. %SIZE WT % (wt %) (ppm) Fe % REC. RECOV. +¼ 20.9 0.57 2.4 24.3 50.6 (15)⅛-¼ 32.3 0.78 2.6 38.8 62.7 (24) 10-⅛  4.89 0.525 3.8 47.3 76.1 (40)16-10 8.49 1.22 3.8 44.3 74.7 (46) 30-16 9.36 1.92 5.8 37.3 84.4 (53)60-30 6.65 pyrite 1.6% 13.56 47.1 sand 5.02% 0.43 75.3 −60 17.3 pyrite2.68% 7.81 69.9 clay 14.62% 1.48 86.5 Au (ppm) = Concentration of goldin size fraction Fe % = Concentration of Fe in size fraction in weightpercent. Au % Rec. = Percent gold recovered from size fraction byperforming a traditional cyanide leach test without biooxidizing orefirst. Bioox. % = Percent gold recovered by cyanide leach after Recov.biooxidation. The percent of biooxidation for each sample is given inparentheses.

After several months of biooxidation, samples were taken from eachcolumn and the percent iron leached noted. The partially biooxidized orewas then leached with cyanide in the same way the original unoxidizedsamples were. The gold extraction of the unoxidized sample and thebiooxidized sample are compared in Table 1. The percent biooxidation foreach size fraction is reported in Table 1 in parentheses. From this dataone can see that the smaller size fractions biooxidized at a fasterrate. Also, all the size fractions show an increase in gold extractionafter being biooxidized.

The +60 to −30 mesh and −60 mesh size fractions were also analyzed forgold extraction. The sand tails from a shaker table separation of therefractory pyrite from the +60 to −30 mesh fraction was fairly low ingold, but the gold was cyanide extractable without biooxidation (75%).The very fine sand and clay from the −60 mesh fraction was higher ingold and in gold extraction (86%). This indicated that no furtheroxidation of the very fine sand and clay materials in this size fractionwas required.

The removal of the small size fractions (i.e., the size fractions havinga particle size less than 30 mesh) including the clay fraction allowedall the columns to have excellent air flow. Columns made with whole oreor whole ore with agglomeration often would become plugged, inhibitingair flow. Thus, by separating the fines and clays, large scale heaps maybe constructed without having to use larger crush sizes (i.e., ¾ inch orlarger) to achieve good air flow.

The pyrite fractions of the −30 and −60 mesh fractions were both high ingold and refractory to cyanide leaching. These pyrite fractions werecombined and then milled to −200 mesh in a ball mill. The −200 meshpyrite concentrate was used in shake flask experiments to determine theamount of gold extraction as a function of percent biooxidation (seeFIG. 4). In preparing these tests, 75 ml of a 500 ppm cyanide solutionwas added to 30 gm of the pyrite concentrate. The solution and ore wasthen rolled at 10 rpm for 96 hrs. before the cyanide solution was testedto determine the amount of gold extracted.

Some of the pyrite from the gravity separated fines was furtherprocessed by grinding to −200 mesh and floating with xanthate to from aconcentrate of over 50% pyrite. A sample of this concentrate weighing500 gm was then mixed with 500 ml of solution containing iron oxidizingbacteria at greater than 10⁸ cells per ml and 3000 ppm ferric sulfate.After one hour, the 500 gm sample of pyrite concentrate suspended in 500ml of ferric-bacteria solution was poured directly onto the top of the+¼ inch ore column, containing about 15 Kg of ore. This was done afterbiooxidation of the column ore had been in progress for over 300 days.The black liquid spread quickly down through the column with most of thepyrite concentrate being retained by the column. The small amount ofpyrite concentrate that did pass through the column was poured back ontothe top of the column and was retained by the column on the second pass.The pyrite appeared to be evenly distributed throughout the column anddid not inhibit the air flow.

Liquid at pH 1.8 was dripped onto the top of the column, as had beendone throughout the experiment. The flow rate was about 200 ml per day.The liquid collected after three days had dropped in Eh from about 650mV to 560 mV. The pH was still at about 1.8 as it had been for a longtime. The iron concentration in the liquid was 2800 ppm, which was justa little lower than the iron concentration of the added bacteriasolution. Two days after adding the pyrite concentrate to the column,the iron concentration in the off solution had increased to 4000 ppm andthe pH had dropped to 1.6 indicating that biooxidation of the pyrite hadstarted. FIGS. 5, 6, and 7 illustrate the change in Eh, pH, and ironconcentration of the column effluent, respectively over time.

Another aspect of the present invention will now be described. In thisaspect, a process for recovering precious metal values from aconcentrate of precious metal bearing refractory sulfide minerals isdescribed. The process comprises (a.) distributing a concentrate ofrefractory sulfide minerals on top of a heap of support material; (b.)biooxidizing the concentrate of refractory sulfide minerals; (c.)leaching precious metal values from the biooxidized refractory sulfideminerals with a lixiviant; and (d.) recovering precious metal valuesfrom the lixiviant.

A concentrate of precious metal bearing refractory sulfide minerals willtypically be prepared from a precious metal bearing refractory sulfideore. The concentrate can be prepared from such ores using well knowngravity separation or flotation techniques. Although gravity separationis cheaper, flotation is the preferred method of separation because ofthe selectivity of the process. The most frequently used collector forconcentrating sulfide minerals in a flotation process is Xanthate.Xanthate flotation processes are well known to those skilled in the artand need not be described in detail herein.

Preferably the particle size of the concentrate is such that 80 to 90%of the concentrate is less than 100 to 300 mesh. More preferably, 80 to90% of the concentrate is less than 100 to 150 mesh.

The optimum size may, however, vary with various ore types. In general,the operator should strive for a particle size which permits optimumseparation in the concentration process and which provides for theoptimal rate of biooxidation versus the incremental costs of additionalfine grinding.

The smaller the particle size of the sulfide minerals within theconcentrate, the more quickly the concentrate will oxidize duringbioleaching. However, the faster biooxidation rate does not alwaysjustify the added energy costs associated with fine grinding an ore or aflotation concentrate.

With the process according to the present invention, the cost of leavingthe concentrate on the heap to biooxidize is minimal. Therefore, aslightly longer biooxidation period may be justified to avoid having toincur additional grinding related expenses. In this regard, the presentprocess has an advantage over mill type processes. In mill typeprocesses, the sulfide mineral concentrate must be very finely ground toensure high biooxidation rates so that the bioreactor can process asmuch concentrate as possible in as short of period of time as possibleto maintain the economics of the process.

After the sulfide mineral concentrate is formed, it is distributed overthe top of a heap of support material. Preferably, the concentrate isdistributed on top of the heap in a slurry form so that the concentratecan be piped directly to the heap without having to be dried first. Thepulp density of the concentrate should be adjusted so that theconcentrate flows well, but does not simply wash through the heap ofsupport material. Because the sulfide mineral particles are hydrophobic,they will tend to stick to the support material rather than migratingcompletely through the heap if the appropriate support material isselected. Nor should blockage of flow channels be a problem if anappropriate size support material is selected.

The purpose of the support material is to capture and retain the sulfideminerals as they slowly migrate down through the heap so that thesupport material acts as a large surface area bioreactor. For thisreason support materials having a high degree of porosity or a roughsurface are preferred since these types of surfaces will tend to captureand retain the concentrate. The more concentrate that the support rockcan support without blockage of the flow channels the better. Supportmaterials that can be used in practicing the present invention includecoarse ore particles, lava rock, gravel, or rock containing smallamounts of mineral carbonate as a source of CO₂ for the biooxidizingbacteria. Lava rock is a particularly preferred support material due toits roughness and high degree of porosity.

Support material which contains a small amount of mineral carbonate isbeneficial not only for the CO₂ that it produces but is also beneficialbecause it will help buffer the acid solution produced as a result ofthe biooxidation process. This will make it easier to control the pH ofthe bioreactor during the biooxidation process.

With respect to selection of an appropriate size of support material,there are several competing interests that should be considered. Smallerdiameter support materials have greater surface area and thus increasethe effective area of the bioreactor created by the heap of supportmaterial. However, smaller diameter support material may be moreexpensive depending on the amount of grinding required to produce thedesired size. Further, smaller diameter support material may be subjectto more blockage of fluid flow channels by the concentrate which isadded to the top of the heap. Larger support material will permit tallerheaps to be formed without risk of flow channels becoming plugged.

Typically, the support material will be larger than about ¼ inch indiameter and smaller than about 1 inch in diameter. Preferably thesupport material is greater than about ⅜ inch in diameter and less thanabout ¾ inch in diameter. A support material having a diameter of about½ inch should be the optimum size.

To biooxidize the concentrate, the heap is inoculated with bacteria orother microbe capable of biooxidizing the sulfide minerals in theconcentrate. Such microbial treatments are well known in the art.Bacteria that can be used for this purpose include Thiobacillusferrooxidans, Leptospirillum ferrooxidans, and Thiobacillus thiooxidans.Thiobacillus ferrooxidans is an especially preferred microorganism forbiooxidation processes.

If the bioleachate solution is recycled, precautionary steps may berequired to prevent toxic materials from building up in the recycledsolution so that the rate of biooxidation is not retarded significantly.The process described in U.S. patent application Ser. No. 08/329,002,filed Oct. 25, 1994, can be used to ensure that inhibitory materials donot build up to the point that they become detrimental to thebiooxidation process.

After the refractory sulfide concentrate is sufficiently biooxidized,the liberated precious metal values can be leached with a lixiviant ofthiourea or cyanide. Cyanide is the preferred lixiviant even though thepH of the heap must first be raised prior to leaching. An advantage ofthiourea is that it is not toxic to the biooxidizing microorganisms. Asa result, intermittent leachings can be performed to dissolve theliberated precious metal values and then the biooxidation process can beresumed.

Dissolved precious metal values can be recovered from the lixiviantusing well known techniques to those skilled in the art such as carbonin leach and carbon in column processes.

Another advantage of the present process is that it can be used as acontinuous process by intermittently adding fresh or new concentrate tothe top of the heap. The advantage of adding fresh concentrate to thetop of the heap is that once the heap is established and biooxidation isoccurring rapidly, the fresh concentrate can be added to maintain thehigh rate of biooxidation within the heap without having to tear downthe heap to process the biooxidized material.

Due to the relatively low capital and operating costs of the presentprocess, it can be used to economically process much lower gradeconcentrates, and as a result lower grade ores, than a mill biooxidationprocess. Further, by distributing the concentrate of precious metalbearing refractory sulfide minerals on top of a heap of supportmaterial, good fluid flow (both air and liquid) is ensured within theheap.

Another aspect of the present invention is now described. In thisaspect, a process is provided for recovering base metal values fromsulfide ores. Such ores include, by way of example, chalcopyrite,sphalorite, nickel sulfide ores, and uranium sulfide ores. The processaccording to this aspect of the invention comprises (a.) forming aconcentrate of metal sulfide minerals; (b.) spreading the concentrate ontop of a heap of support material; (c.) biooxidizing the concentrate;and (d.) recovering metal values from the solution used to biooxidizethe metal sulfide minerals. Due to the fact that this process, like theprocess previously described for processing concentrates of preciousmetal bearing sulfide minerals, uses a heap of support material for thebioreactor, its capital and operating costs are less than that of a millbioleaching operation. However, due to the good air flow in the heap,the biooxidation rate of the sulfide minerals is quite high and canapproach that of what would be observed in a mill type operation.

Depending on the sulfide ore from which the concentrate is obtained, thebase metal values that can be recovered from the process according tothe present aspect of the invention include copper, zinc, nickel anduranium.

The process parameters and considerations for the process according tothe present aspect are much the same as those set forth above for themethod of processing precious metal bearing concentrates of refractorysulfide minerals. The primary difference between the two processes,however, is that the base metal values of interest in the presentprocess dissolve during the biooxidation process. As a result, the metalvalues are recoverable directly from the solution used to biooxidize theconcentrate of metal sulfide minerals. The technique used to extract themetal values of interest will depend on the specific metal of interest.As those skilled in the art will immediately recognize, such techniquesmay include solvent extraction, iron cementation, and precipitationthrough pH adjustments. Solvent extraction is a particularly preferredmethod of removing copper from the bioleachate solution.

As with the above described process for recovering precious metal valuesfrom a precious metal bearing concentrate of sulfide minerals, thepresent process can be operated in a continuous mode by addingconcentrate on an intermittent basis. For example, concentrate can beadded on a daily or weekly basis. As described above, such additionswill ensure that the rate of biooxidation remains high for theconcentrate that is distributed over the heap and which has migratedthrough the heap.

As one skilled in the art will recognize, the process according to thepresent aspect of the invention can be combined with the above processfor recovering precious metal values from a concentrate of refractorysulfide minerals. This is because base metal values from the refractorysulfide minerals will inherently dissolve into the bioleachate solutionduring the biooxidation process while simultaneously liberating anyoccluded precious metal values in the sulfide minerals. These values canthen be recovered if desired using the techniques described above.

EXAMPLE 3

Two simultaneous bioleaching tests were set up to test the rate ofbiooxidation of a gold bearing ore pyrite concentrate. The first testconsisted of a column type experiment to simulate a heap leachingprocess and the second consisted of a shake flask experiment to simulatea stirred tank process.

The starting concentrate for both tests was obtained from the Jamestownmine in Tuolumne County, Calif. The mine is owned by Sonora Goldcorporation and lies along the mother lode vein system. The concentratewas produced using a xanthate flotation process and contained 39.8%sulfides and 36.6% iron. The sulfide minerals within the concentrateprimarily consisted of pyrite. Size analysis showed that over 76% of theconcentrate particles were smaller than 200 mesh. The concentrate had ahigh gold concentration (about 2 oz. per ton of concentrate) and wasknown to be refractory to cyanide leaching.

The percentage of biooxidation in each of the tests was determined byanalysis of the iron concentration in all solutions removed from thecolumn or in the case of the flask experiment the concentration of ironin solution plus any iron solution removed.

A culture of Thiobacillus ferrooxidans was used to biooxidize thesulfide mineral concentrate in each of the tests. The culture ofThiobacillus ferrooxidans was originally started with ATCC strains 19859and 33020. The culture was grown in an acidic nutrient solution having apH of 1.7 to 1.9 and containing 5 g/l ammonium sulfate ((NH₄)₂SO₄)),0.833 g/1 magnesium sulfate heptahydrate (MgSO₄.7H₂O), and 20 g/l ironin the form of ferrous and ferric sulfate. The pH of the solution wasadjusted to the above range using sulfuric acid (H₂SO₄).

Prior to application of the culture to the test samples, the mixedculture of sulfide mineral oxidizing bacteria was grown to a celldensity of 4×10⁹ to 1×10¹⁰ cell per ml.

The column experiment was started by inoculating a 150 g sample ofconcentrate with about 10⁸ cells per gram of concentrate. This wasaccomplished by adding three milliliters of bacteria at 5×10⁹ cells permilliliter to the 150 g sample of pyrite concentrate. The 150 g ofpyrite concentrate suspension was then poured into a 3 inch by 6 footcolumn filled about halfway with 3 liters of ⅜ inch lava rock. The lavarock support material was chosen because it has a high surface area andit holds up well to the acid condition encountered during biooxidation.

During inoculation and subsequent solution additions, the pyriteconcentrate did not wash out of the column. Most of the pyriteconcentrate was held in the first foot of the lava rock. Air and liquidwere introduced through the top of the column. The bioleach solution wasrecirculated until the pH of the column was adjusted down to about 1.8.After biooxidation started within the column, a 0.2×9K salts solutionhaving a pH of 1.8 and containing 2000 ppm of iron, primarily in theferric form, was fed to the column. The 2,000 ppm of iron was subtractedfrom all analysis of iron in solution coming off of the column.

The composition of the standard 9K salts medium for T. ferrooxidans islisted below. The concentrations are provided in grams/liter.

(NH₄)SO₄ 5 KCl 0.17 K₂HPO₄ 0.083 MgSO₄•7H₂O 0.833 Ca(NO₃)•4H₂O 0.024

The notation 0.2×9K salts indicates that the 9K salt solution strengthwas at twenty percent that of the standard 9K salts medium.

After 26 days of biooxidation, about 35% of the iron in the pyriteconcentrate had been oxidized. At this point, the test was converted toa continuous process test by adding 3 g of new concentrate to the columnevery day. After 9 more days, the rate of pyrite addition was increasedto 6 g per day.

The flask experiment was started at the same time as the columnexperiment. To start the experiment, a 50 g sample of the pyriteconcentrate was inoculated with 1 milliliter of the bacteria culture.The pyrite concentrate was then added to 1 liter of 0.2×9K saltssolution having a pH of 1.8 in a large shake flask. Not only was theconcentrate inoculated with the same bacteria, but it was alsoinoculated at the same number of cells per gram.

Air was introduced to the bioleach solution by orbital shaking of theflask at about 250 rpm. Solution was removed from the flask from time totime to keep the ferric concentration from getting much higher than thatin the column.

When the column experiment was converted to a continuous process on day26, the flask experiment was also converted to a continuous test byadding 1 g of pyrite concentrate per day to the flask. After 9 moredays, the amount of concentrate added was increased to 2 g per day.

After 58 days, the pyrite additions to both the flask and columnexperiments were stopped. Both the column and the flask were thenallowed to biooxidize an additional 20 days. At this point, theconcentrate in the column was about 76% oxidized and the concentrate inthe flask was about 89% oxidized. The column was then leached for 10days with thiourea to extract liberated gold. The thiourea onlyextracted about 30% of the gold. However, after 3 days of reverting backto additions of the 0.2×9K salts solution having a pH of 1.8 andcontaining 2,000 ppm of ferric iron, the Eh and the iron concentrationof the column effluent increased. This indicated that the thiourea wasnot toxic to the bacteria and that thiourea extractions could be donefrom time to time without killing the bacteria.

FIG. 8 shows the amount of biooxidation versus time in days for both thecolumn and flask concentrate bioleaching tests. The phrase “TU leach” inFIG. 8 stands for thiourea leach. The data used to prepare FIG. 8 istabulated in Tables 2 and 3 at the end of this example.

As indicated above, the flask was meant to simulate a stirred tankprocess. When the flask test was converted to a continuous process byadding pyrite each day, it was meant to simulate a large scale processin which new pyrite is introduced on an intermittent basis to a rapidlybiooxidizing tank containing a large amount of bacteria that haveadapted to the ore. The daily addition of pyrite to the column was doneto test the feasibility of a continuous process in which a concentrateof precious metal bearing sulfide minerals is continuously orintermittently added to the top of a heap comprised of biooxidizingconcentrate distributed on a heap of support material such as lava rock.

As the above tests demonstrate, the rates of biooxidation were notsignificantly different between the column and flask tests. The start ofbiooxidation was a little slower in the column test. This may have beendue to about a 10 day lag time in adjusting the pH of the column down to1.8. The rate of biooxidation in the column then picked up to be thesame as the flask. Later in the experiment the rate began to slow downagain. This may have been due to a lack of mixing of the fresh pyritewith the biooxidizing pyrite. However, the rates of biooxidation betweenthe two tests were close enough to demonstrate the viability of theprocess according to the present invention. The viability of the presentprocess is especially attractive in view of the much lower capital andoperating costs of a heap process as compared to a stirred tank process.

TABLE 2 Data From Column Biooxidation Test G. of iron Total g. of %bioox. of Time in added to iron pyrite based Conc. of days columnremoved on iron iron in g./l 0 54.400 2.830 5.200 1.884 15 54.400 5.50010.100 2.840 20 54.400 12.617 23.180 4.704 23 54.400 14.480 26.620 4.97626 55.540 19.230 34.620 9.088 27 56.630 20.430 36.070 9.432 28 57.72022.329 38.700 9.800 29 58.800 23.987 40.800 6.400 30 59.900 25.17642.000 5.964 31 61.000 27.075 44.380 5.876 32 62.070 28.337 45.650 6.50833 63.160 29.285 46.360 6.212 34 64.250 30.257 47.080 4.900 35 65.34031.824 48.700 7.224 36 69.700 32.970 47.300 5.428 37 69.700 34.06648.900 5.265 38 74.050 35.184 47.500 5.620 39 74.050 36.302 49.000 4076.230 37.420 49.100 5.120 41 78.410 38.425 49.000 5.000 42 80.59039.453 48.900 5.024 43 82.760 40.744 49.200 5.536 44 84.940 42.17249.600 5.808 45 89.300 43.602 48.800 5.964 46 89.300 45.032 50.400 4791.480 46.462 50.800 5.976 48 93.660 47.932 51.180 6.200 49 95.83649.650 51.800 6.896 50 98.014 50.582 51.600 7.328 51 100.192 52.14252.040 8.240 53 104.548 55.591 53.170 9.664 54 106.726 58.012 54.3608.052 55 108.896 59.835 64.950 8.288 56 111.066 61.571 55.440 8.200 57113.236 63.136 55.760 7.304 58 115.406 65.370 56.640 8.384 59 115.40667.640 58.610 8.484 61 115.406 70.806 61.350 8.208 62 115.400 72.34462.690 7.128 63 115.400 72.777 63.930 6.776 64 115.400 75.013 65.0005.852 65 115.400 76.169 66.000 5.728 66 115.406 77.325 67.000 5.728 68115.406 79.668 69.030 5.748 69 115.406 80.468 69.730 4.668 70 115.40081.043 70.220 4.740 71 115.400 81.828 70.904 4.856 72 115.400 82.71671.674 5.064 73 115.400 83.781 72.590 4.804 75 115.400 84.975 73.6304.488 76 115.400 85.609 74.180 4.112 90 115.400 87.170 75.533 2.892 93115.400 88.754 76.900 3.476

TABLE 3 Flask Biooxidation Data Flask % Time in days bioox. by iron 04.930 15 18.890 21 29.850 26 37.400 28 39.790 36 45.370 44 46.890 4752.310 49 54.510 51 58.380 54 62.010 56 62.630 58 65.400 63 72.110 6981.410 72 83.300 75 89.440

Although the invention has been described with reference to preferredembodiments and specific examples, it will readily be appreciated bythose of ordinary skill in the art that many modifications andadaptations of the invention are possible without departure from thespirit and scope of the invention as claimed hereinafter. For example,while some of the processes according to the present invention have beendescribed in terms of recovering gold from refractory sulfide orrefractory carbonaceous sulfide ores, the processes are equallyapplicable to other precious metals found in these ores such as silverand platinum.

1. A method for recovering metal values from a sulfide ore, the methodcomprising: a. crushing the refractory sulfide ore to a size in therange of approximately 0.25 inch to 1 inch; b. separating the crushedrefractory sulfide ore into a fines fraction and a coarse fraction; c.forming a heap with the coarse fraction of the refractory sulfide ore;d. producing a concentrate of refractory sulfide minerals from the finesfraction; e. adding the concentrate to the heap; f. biooxidizing the orein heap, including the concentrate of sulfide mineral particles; and g.leaching precious metal values from the biooxidized sulfide mineralparticles.
 2. The method of claim 1, wherein the precious metal valuesinclude gold.
 3. The method of claim 1, wherein the crushed ore isseparated into the fines fraction and the coarse fraction by wetscreening.
 4. A method for recovering metal values from a sulfide ore,the method comprising: a. crushing the refractory sulfide ore to a sizein the range of approximately 0.25 inch to 1 inch; b. separating thecrushed refractory sulfide ore into a fines fraction and a coarsefraction; c. forming a heap with the coarse fraction of the refractorysulfide ore; d. producing a concentrate of refractory sulfide mineralsfrom the fines fraction; e. adding the concentrate to the heap; f.biooxidizing the ore in heap, including the concentrate of sulfidemineral particles; g. leaching metal values from the biooxidized ore;and h. recovering metal values leached from the ore.
 5. The method ofclaim 4, further comprising treating the separated fines to recovermetal values contained therein.
 6. The method of claim 4, wherein therecovered metal is copper.
 7. The method of claim 4, wherein therecovered metal is at least one selected from the group consisting ofgold, silver, and platinum.
 8. The method of claim 7, wherein therecovered metal is gold.
 9. The method of claim 4, wherein the crushedore is separated into the fines fraction and the coarse fraction by wetscreening.
 10. A method for recovering metal values from a concentrateof sulfide mineral particles prepared from refractory sulfide ore, themethod comprising: a. adding the concentrate of sulfide mineralparticles to a heap of support material having a partial size in therange of approximately 0.25 inch to 1.0 inch; b. biooxidizing theconcentrate of sulfide mineral particles; c. leaching metal values fromthe biooxidized sulfide mineral particles; and d. recovering metalvalues leached from the biooxidized sulfide mineral particles.